Method for nickel and cobalt recovery from laterite ores by reaction with concentrated acid and water leaching

ABSTRACT

A process for leaching laterite ores containing limonite and saprolite in a two stage process. The first stage consisting of mixing and reacting the ore with concentrated mineral acid, and the second stage consisting of preparing a slurry of the acid/ore mixture in water and leaching the mixture to dissolve nickel and cobalt. Iron is efficiently separated from nickel and cobalt in the solid leach residue primarily as an oxide or hydroxide of ferric iron other than jarosite.

CROSS REFERENCE TO RELATED APPLICATION

This application claims the benefit of U.S. provisional patentapplication No. 60/583243 filed Jun. 28, 2004, the disclosure of whichis hereby incorporated herein by reference.

FIELD OF THE INVENTION

The present invention relates to the hydrometallurgical processing ofnickeliferous laterite ore, and in particular to a method for acidleaching both the limonite fraction and the saprolite fraction of suchores in a single process.

BACKGROUND OF THE INVENTION

Laterite ores are formed by the in-situ weathering of nickel-bearingultramafic rocks near or at the surface of the earth in tropicalenvironments by the action of naturally acidic meteoric waters overgeologic time. They consist of a variety of clay, oxide and silicateminerals, some enriched in nickel and/or cobalt, and this distinguishesthem from the other major class of nickel ores, the sulfide ores. Thelatter consist typically of sulfide minerals of iron, nickel and cobalt,often with copper and minor precious metals, and are associated withmafic-ultramafic magmatic intrusions in the earth's crust.

The weathering process typically creates a layered deposit, with theproducts of complete or most extensive weathering occurring near thesurface and these grading into the products of lesser degrees ofweathering as depth is increased and finally terminating in unweatheredrock at some greater depth. The highly weathered layer usually containsmost of its contained nickel microscopically distributed within veryfinely divided goethite particles. Goethite is an oxyhydroxide of ferriciron with the chemical formula FeOOH. This layer is usually given thename limonite, and typically contains a high proportion of iron.

Cobalt is usually associated with the limonite layer and is usuallypredominantly associated with oxidized manganese minerals (Mn(III)and/or Mn(IV) containing oxides and hydroxides), often called asbolaneor manganese wad.

The lesser weathered layers typically contain increasing proportions oftheir contained nickel in various magnesium silicate minerals, such as,for example, serpentine. Serpentine is a silicate mineral of magnesiumwhich has the chemical formula 3MgO*2SiO₂*2H₂O. It is believed thatnickel substitutes for some of the magnesium in serpentine. Magnesiummay also be substituted by other divalent metals, for example, ferrousiron (Fe²⁺). There may be many other silicate minerals that also hostnickel in the incompletely weathered zones. The partially weathered,high-magnesium bearing zone is often given the name saprolite, orgarnierite. (“Garnierite” is also used to describe a particularapple-green colored magnesium-nickel silicate mineral of variablecomposition.)

In some deposits there is another zone typically located between thelimonite and saprolite that consists predominantly of nontronite clays;these are silicates of magnesium, iron and aluminum that may also benickeliferous. In most deposits located in the (current) tropics, thenontronite zone is largely absent.

It should be noted also that none of the weathering zones arehomogeneous in mineralogical or chemical composition, nor is theboundary between the zones parallel to the earth's surface. However,there is usually a fairly sharp transition from ore of high iron andrelatively low magnesium contents to ore of a relatively high magnesiumcontent and lower, although variable, iron content, which occurs oververtical distances of 1 to 3 m within a laterite deposit.

For illustration purposes only, typical chemical compositions oflimonite and saprolite are as follows: Limonite: 1.0-1.8% Ni, 0.05-0.3%Co, 35-50% Fe, 0.2-3.5% Mg Saprolite: 1.2-3.5% Ni, 0.02-0.07% Co, 7-20%Fe, 10-20% Mg

Each zone also contains typically significant concentrations ofaluminum, manganese and chromium, as well as trace concentrations ofother heavy metals such as copper and zinc in a variety of otherminerals.

A challenging aspect of nickel recovery from laterite ores is that thenickel values typically can not be concentrated substantially byphysical means, that is, so-called ore dressing techniques, prior tochemical processing to separate the metal values. This renders theprocessing of laterites expensive, and means to lower the costs ofprocessing laterites have been sought for many decades.

Also, because of the distinct mineralogical and chemical composition oflimonite and saprolite ores, these ores usually are not amenable toprocessing by the same process technique.

One known acid leaching process for nickel laterites is the so-calledHigh Pressure Acid Leaching (HPAL) process (see, for example pages437-453 in “The Winning of Nickel Its Geology, Mining and ExtractiveMetallurgy,” by Joseph R. Boldt, Longmans Canada Ltd. 1967). Thisprocess was first employed at Moa Bay in Cuba in the late 1950s andadditional plants were constructed in Western Australia in the late1990s.

The process utilizes sulfuric acid leaching at high temperature,typically 250° C., and high pressure; the associated steam pressure at250° C. is approx. 570 psi. At this temperature, the nickeliferousminerals in the ore are nearly completely solubilized. The dissolvediron is rapidly precipitated as hematite (Fe₂O₃) at the high temperatureemployed because this compound is largely insoluble even in slightlyacidic solutions at this temperature. The nickel remains in solution andafter cooling, the leach residue containing iron is separated from thenickel-bearing solution by thickening in a series of wash thickeners, aso-called counter-current decantation (CCD) circuit. Thus, the primaryobjective of the leaching process, which is separation of nickel fromiron, is achieved.

A major disadvantage of the HPAL process is that it requiressophisticated high-temperature autoclaves and associated equipment whichare expensive, both to install and to maintain. In addition, the HPALprocess also consumes more sulfuric acid than is required tostoichiometrically dissolve the non-ferrous metals content of the orebecause at high temperature most of the sulfate ions provided bysulfuric acid are tied up as bisulfate ions (HSO₄ ⁻). In other words,sulfuric acid (H₂SO₄) only dissociates to release a single proton (H⁺)at high temperature. On cooling and neutralization of the leach liquorthe bisulfate ions decompose to sulfate (SO₄ ² ⁻) and another proton.The latter proton (acid) is therefore not utilized fully for leachingand results in excess sulfuric acid which must be neutralized, forexample with limestone.

Another disadvantage of the HPAL process is that it is limited totreating largely limonite-type feeds because the presence of saprolitewill cause a large, and often uneconomic, increase in sulfuric acidconsumption due to the leaching of magnesium from saprolite. This isexacerbated by the bisulfate “shift” problem at high temperature, whichis described above.

U.S. Pat. No. 4,097,575 describes an improvement to the HPAL processwhich constitutes roasting saprolite ore below about 820° C. in order torender the ore more reactive with sulfuric acid and then using theroaster calcine to neutralize excess acid in the discharge of anautoclave wherein pressure leaching of limonite ore occurs. Nickelcontained in the saprolite ore is largely dissolved during thisneutralization. The reported advantages of this process are that itbetter utilizes the sulfuric acid added during pressure leaching oflimonite, it reduces the consumption of limestone or other costlyneutralizing agents to treat the autoclave discharge liquor, and itachieves the capability of treating both the limonite and saprolitefractions of a typical nickel laterite orebody. Disadvantages of theprocess are that it still requires the use of expensive autoclaves forleaching limonite, and it requires a roasting process for saprolite ore,which is expensive both in capital and operating cost terms.

U.S. Pat. No. 6,379,636 B2 describes a further improvement to theprocess described in U.S. Pat. No. 4,097,575 wherein the saproliteroasting step is eliminated and the saprolite in “raw” form is used toneutralize the excess acid in the autoclave discharge solution. Inaddition, more acid could be added to the discharge to increase theamount of saprolite that could be leached. However, this process stillrequires the use of expensive autoclaves.

Several processes have also been described that utilize acid leaching atatmospheric pressure only, eliminating the disadvantages of pressureleaching described above. U.S. Pat. No. 3,793,432 describes anatmospheric leaching process for laterite ore, in which the ore isreacted with sulfuric acid at or below the boiling point and theprecipitation of dissolved iron is achieved by the addition of an ironprecipitating agent such as ammonium, sodium, potassium or lithium ions.Although not stated explicitly, all of the examples cited in thespecification employed limonite ore samples, as evidenced by the highiron content and low magnesium content of the feed ore. While thisprocess overcomes the disadvantages of pressure leaching, it has otherdisadvantages. First, the precipitation of iron would be as a jarositecompound, which is a thermodynamically unstable compound of iron thatdecomposes over time to release sulfuric acid, thus causingenvironmental problems. (Although jarosite is not stated explicitly itwould be apparent to one skilled in the art that jarosite willprecipitate at the conditions outlined in the examples.) Jarositecontains two moles of sulfate for every three moles of iron and thusthis compound represents substantial excess consumption of sulfuric acidto provide the necessary sulfate ions.

Second, the nickel extractions from the ore were apparently relativelylow. While extractions were not stated explicitly, based on the nickelcontent of the residue and the fact that the residue weight must be morethan the weight of the original ore because jarosite contains a lowerpercentage of iron than the original ore and virtually all of the ironremained in the residue, nickel extractions were usually in the 60-65%range. Third, there is a requirement for very long leach times, of theorder of 4-5 days. Fourth, there is a need to add relatively expensiveiron precipitating agents such as potassium carbonate, sodium carbonateor the like.

U.S. Pat. Nos. 6,261,527 B1 and 6,680,035 B2 describe anotheratmospheric leaching process in which limonite ore is first “totally”leached with strong sulfuric acid, i.e. both nickel and iron aresubstantially dissolved from goethite, and then saprolite ore is leachedin the resulting limonite leach slurry while simultaneouslyprecipitating iron as jarosite by the addition of a jarositeprecipitating agent. This process also has the disadvantages ofproducing jarosite, requiring separate mining and preparation of thelimonite and saprolite fractions of the ore, and being limited to anarrow range of saprolite to limonite ratios. The latter disadvantage isdue to the fact that the quantity of saprolite that can be leachedeffectively is determined by the quantity of ferric iron in the limoniteleach solution.

WO 03/093517 A1 describes an improvement to this process, whichconstitutes eliminating the addition of a jarosite-forming ion such assodium, potassium and ammonium, and causing the iron to precipitate as acompound other than jarosite, such as goethite. The process overcomesthe disadvantages of jarosite, but sulfuric acid consumption was 0.59 to0.87 tonnes per tonne of ore in the examples cited, and was over 0.72tonnes per tonne of ore in nine of the eleven examples cited.

The processes described in U.S. Pat. Nos. 6,261,527 B1 and 6,680,035 B2and WO 03/093517 A1 are based on the fact that goethite is morerefractory to acid leaching than typical saprolite minerals, such asserpentine. This has been demonstrated by other researchers (see, forexample: John H. Canterford, “Leaching of Some Australian NickeliferousLaterites with Sulfuric Acid at Atmospheric Pressure,” Proc.Australasian Inst. Min. Metall., 265 (1978), 19-26; N. M. Rice and L. W.Strong, “The Leaching of Lateritic Nickel Ores in Hydrochloric Acid,”Canadian Metallurgical Quarterly, 13(3)(1974), 485-493; and FIG. 5 ofU.S. Pat. No. 5,571,308). Thus, only saprolite can be used effectivelyin the second stage of leaching where iron precipitation occurssimultaneously. This is because the acidity of the solution must berelatively low to enable the precipitation of jarosite and even lower toenable the precipitation of goethite or other hydrolysis products offerric iron. The goethite contained in limonite would leach very slowlyunder these conditions. Limonite (principally goethite) is thus leachedin an initial stage with a relatively high acid concentration andboth-iron and nickel are brought into solution.

U.S. Pat. No. 3,244,513 describes a process in which laterite ore,largely of limonitic type (defined as >25% iron), is mixed withconcentrated sulfuric acid in the presence of limited water, then themixture is roasted at temperatures from approx. 500 to 725° C. in orderto cause preferential sulfation of the nickel, cobalt, magnesium andmanganese values over iron. Subsequent water leaching results in highextraction of nickel and cobalt and low extraction of iron to solution.The advantages of this process are that it does not require expensiveautoclaves to effect leaching. The primary disadvantage is that itrequires an expensive roasting step.

U.S. Pat. No. 4,125,588 describes a similar process as that described inU.S. Pat. No. 3,244,513, except that the roasting step is omitted andthe mixing of ore with concentrated acid is done in a carefullycontrolled fashion wherein ore is first mixed with concentrated sulfuricacid in the absence of significant water and then water is added incontrolled amounts to initiate sulfation of the ore, and then finallythe mixture is leached with further amounts of water. The advantage ofthis process is that it eliminates the roasting step required in U.S.Pat. No. 3,244,513. However, the process also has significantdisadvantages.

One disadvantage is that the ore used should contain no more than 1%moisture, which means that in most cases the ore must be dried, asin-situ moisture contents of laterite ores are usually 20% or more. Asecond disadvantage is that the process does not provide selectivedissolution of nickel over iron, as illustrated in all of the examplescited in the patent (>90% iron extraction). The separation of nickelfrom iron in solution usually results in additional nickel losses.

In addition, this process is only applicable to ores containing “largeamounts of magnesia and silica,” i.e. the saprolitic or garnieriticores. While the precise ore composition is not cited in all of theexamples, it is evident from the data given that the ore contained from3 to 4 times as much magnesium as iron, clearly indicating thatlimonitic ores (iron/magnesium ratios of from approx. 10 to 90 byweight) were not considered. Moreover, acid consumption is very high inthis process, from about 0.9 to 1.1 tonnes of sulfuric acid per tonne ofore.

U.S. Pat. No. 3,093,559 describes another process for treatment oflateritic ores with relatively concentrated sulfuric acid (from about 25to 50% sulfuric acid). In this process sufficient acid is added to causesulfation of most or all of the metal values including iron. Iron isthen separated from nickel by evaporating the leach solution to drynessand roasting the resulting salts at 975 to 1050° F. in order to convertthe iron to hematite. Subsequent releaching of the calcine brings nickelinto solution, leaving iron in the residue. As with various processesdescribed above, the requirement for a high temperature roasting step isa significant disadvantage in this process.

U.S. Pat. No. 2,899,300 describes a process in which moist laterite oreis treated with concentrated sulfuric acid and the mixture is then driedby baking at a temperature between 100° C. and 150° C., preferably 125°C., before water leaching to dissolve the metal values in the ore. Thebaking step is a significant disadvantage because it requiressubstantial heat energy to evaporate the contained water in the ore/acidmixture. Furthermore, as illustrated by the example given in thespecification, nickel dissolution is relatively low, while irondissolution is relatively high, ˜77% and 53%, respectively. Nickelextraction can reportedly be increased by adding additional sulfuricacid to the first stage residue and carrying out a second sulfation andwater leaching step, but this adds complexity to the process and willnot improve the iron/nickel separation achieved in the process.

U.S. Pat. No. RE37,251 describes a process for the pressure leaching ofnon-cuprous ores and concentrates, including nickel laterite ores, usingan acidic solution of bisulfate and sulfate ions along with halogenions, e.g. chloride ions, and oxygen. According to the specification,the temperature and pressure required are 225° C. and 450 psig O₂. Giventhat the steam pressure at 225° C. is approximately 370 psi, the totalpressure would be in the range of 820 psig. These conditions are fairlysimilar to the high pressure acid leach conditions and would thereforerequire the use of expensive high pressure autoclave systems, asdescribed above.

The object of the present invention is to obviate or mitigate thedisadvantages of these known processes. It is a further object toprovide a process to acid leach a mixture of limonite and saproliteores, such as would be produced by bulk mining of a typical lateriteorebody without any subsequent separation of ore types, underatmospheric or low pressure conditions, while achieving high extractionsof nickel and cobalt and very low ultimate extraction of iron.

BRIEF SUMMARY OF THE INVENTION

The present invention provides a process for the efficient leaching ofnickel and cobalt from limonite and saprolite ore, such as could beproduced by bulk mining of a typical nickel laterite deposit, in a twostage process, the first stage consisting of mixing and reacting the orewith concentrated mineral acid, and the second stage consisting ofpreparing a slurry of the acid/ore mixture in water and leaching themixture to dissolve nickel and cobalt. Iron is efficiently separatedfrom nickel and cobalt in the solid leach residue primarily as an oxideor hydroxide of ferric iron other than jarosite.

The process may also include curing of the mixed ore and acid prior towater leaching, and the ore is first crushed before it is mixed andreacted with the acid. Preferably, the saprolite is crushed separatelyand then blended with the limonite. More preferably, the acid is mixedfirst with the limonite, and the saprolite is added subsequently.

The water leaching is advantageously carried out at a modestly elevatedtemperature. For atmospheric leaching, the temperature is preferablywithin the range 95-105° C. Alternatively, faster leach times may beobtained by leaching in an autoclave with a temperature up to about 150°C. The corresponding pressure is in the range of up to about 70 psia andis approximately equal to the saturated steam pressure at the leachingtemperature.

The acid used is preferably selected from sulfuric, hydrochloric, ornitric acid, and is more preferably sulfuric acid. A mixture of any ofthese acids may also be used.

Advantageously, a reductant, such as sulfur dioxide, hydrogen sulfide,soluble bisulfite and sulfite compounds, or soluble ferrous ironcompounds, is added during the water leaching to enhance dissolution ofcobalt.

The process of the present invention avoids the high cost ofhigh-pressure autoclaves and reduces the production of jarositecompounds. In some embodiments, it also avoids the need to separatelymine and separately treat limonite and saprolite ore types, and it mayallow processing of ore over a wide range of saprolite to limoniteratios.

It has been found that the present invention can achieve at least about80% nickel extraction and as much as 95% or more cobalt extraction, withless than about 15% iron extraction.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a flow sheet showing in simplified form one embodiment of theprocess of the present invention.

FIG. 2 shows another embodiment of the process of the present inventionin which some of the leach residue is recycled in order to provide seedfor iron precipitation.

FIG. 3 shows a third embodiment of the process of the present inventionin which all of the required acid is first mixed with the limonite oreand then the saprolite ore is mixed with the resulting limonite/acidmixture.

FIG. 4 shows a fourth embodiment of the process of the present inventionin which all of the required acid is first mixed with the limonite ore,and then crushed saprolite ore and water are added to the resultinglimonite/acid mixture, after which the resulting mixture is ground andthen subjected to atmospheric leaching.

DETAILED DESCRIPTION OF THE INVENTION

The present invention provides an improved process for the extraction ofnickel and cobalt from nickeliferous laterite ore, while rejecting mostof the iron contained in the ore in a solid leach residue. The processdoes not require the prior separation of the limonite ore from thesaprolite ore and can treat a mixture of the two ore types. However,where it is convenient to mine limonite and saprolite ore separately,high nickel extraction can be achieved by first reacting acid directlywith the limonite ore and then adding the saprolite ore.

Referring to FIG. 1, run-of-mine laterite ore, consisting of a mixtureof limonite and saprolite, is first crushed to approximately 5 to 10 mmtop size. The crushed ore is then mixed with a concentrated mineralacid, chosen from the group sulfuric, hydrochloric and nitric, or amixture of these, in a suitable device such as a pug mill. It is notnecessary to dry the ore, which typically contains substantial freemoisture in the run-of-mine condition, prior to the addition of theconcentrated acid.

The quantity of acid added is at least that required tostoichiometrically dissolve the soluble non-ferrous metals in the ore(but not the iron), i.e. most of the nickel, cobalt, magnesium,aluminum, copper, zinc, and a small portion of the chromium, in the oreduring the subsequent leaching stage. A small excess of acid is added toprovide some free acidity in the subsequent leaching stage, so as toleach a small proportion of the iron, and to ensure maximum extractionof nickel and cobalt. Acid addition is limited to ensure that final ironextraction from the ore slurry is minimized. In some cases, some of themagnesium and aluminum may be insoluble and this should be taken intoaccount to determine the precise acid addition.

Water addition during the acid mixing process is minimized to the extentpossible to provide the highest concentration of acid for reaction withthe ore minerals. Water addition is only necessary if the ore/acidmixture stiffens so much that it can not be blended completely or easilyhandled otherwise. The addition of concentrated acid to moist oreresults in the generation of substantial heat, raising the temperatureof the mixture as high as, and even above, the boiling point of waterand causing significant water evaporation. Additional water can be addedduring the blending process, if desired, to control the consistency ofthe acid/ore mixture to that of a fluid paste. If no water is added,depending on the moisture content of the ore and the net amount of heatgenerated, it may also be possible to form a dry, powdery reactionproduct. (Intermediate between the conditions of a fluid paste and adry, powdery material, a stiff, toffee-like mixture may be formed, whichmay be more difficult to handle.)

Ideally, the mixture will have either a fluid-paste consistency or adry, powdery consistency to permit easy handling. The resulting acid/oreblend, or “pug”, may be allowed to “cure” at ambient temperature for asufficient time as to allow substantial reaction between the mineralacid and the mineral constituents of the ore. This can be done bystockpiling the pug on a prepared impermeable pad and allowing the pugto stay on the pad for as much as several days before reclaiming thematerial and subjecting it to the next phase of the process.

To facilitate stockpiling and reclaiming, it may be desirable to formsmall, discrete portions of the pug by, for example, extruding orpelletizing the material before stockpiling. However, it has been foundthat satisfactory results may be achieved with minimum or no curingtime. In this case, the acid/ore mixture may be directly blended withwater. Subsequent grinding may not be necessary if the ore was groundprior to adding acid. Longer curing may give slightly better nickelextraction and lower iron extraction in the subsequent water leachingstep.

After curing, the pug is ground to a particle size sufficient forleaching in agitated tanks. Conveniently, the water required to form aleach slurry can be added prior to grinding so that this step can becarried out in the wet mode. The pug should be ground to a particle sizesufficient to enable off-bottom solids suspension of most of theparticles in the slurry without requiring excessive agitation powerinput. It is desirable to make the slurry as dense as possibleconsistent with good mixing during leaching in order to minimize waterrequirements and produce the most concentrated nickel and cobalt-bearingpregnant leach solution.

The resulting leach slurry is heated, as required, in a batch leachreactor or series of leach reactors if the process is to be carried outcontinuously. The temperature of the leach slurry is preferablymaintained at or near the normal boiling point of the leach liquor,typically 95-100° C. Live steam injection, or other means of addingthermal energy, can be used for this purpose. Some of the heat requiredis provided by the heat of solution of excess acid and metal saltsformed during the acid blending step. In addition, further amounts ofheat may be recovered during the pugging process itself by cooling thepugging reactor with the water to be used in leaching, thus preheatingthe water, or by recovering steam generated during the acid orereaction.

The leach mixture is allowed to react until most of the nickel andcobalt in the ore have dissolved and most of the iron is present in thesolid leach residue. The agitated leach retention time is of the orderof 12 to 48 hours. Satisfactory results are typically obtained in about15 hours or less. In some cases, it may be desirable to add a reducingagent in controlled amounts to the leach slurry in order to enhance theleaching of cobalt associated with oxidized manganese minerals, e.g.asbolane, in the ore. For example, controlled additions of sulfurdioxide can be added to improve cobalt (and manganese leaching), as willbe apparent to those skilled in the art. Other reductants capable ofreducing trivalent and tetravalent manganese oxides to the divalentstate may also be used.

The leaching process can be accelerated significantly by employing atemperature above the boiling point. For example, by heating the leachslurry to about 150° C. in an autoclave, the leach time can be reducedto about 1 hour duration. Further increases in temperature may bebeneficial but it is an aim of this invention to eliminate thecomplexity associated with high pressure leaching. It should be notedthat the pressure required to leach the pug plus water slurry at 150° C.is only about 70 psia (the steam pressure at 150° C.). At this pressure,the slurry can be pumped to the autoclave with conventional centrifugalpumps and the pressure letdown system can be a very simple single stagevalve or choke. Suitable autoclave equipment to be used in the presentinvention is vastly different from that required for high pressureleaching at temperatures of around 250° C. and pressures of around450-500 psia, as will be apparent to those skilled in the art.

The use of a temperature above the boiling point may also provide ahigher nickel/iron ratio in solution, which is advantageous with respectto downstream processing of the leach solution. This is because in mostcases virtually all iron must be removed from solution before effectingnickel and cobalt recovery. Usually, the residual iron in solution isremoved by adding a base, for example calcium carbonate, to the leachslurry and precipitating iron oxyhydroxide compounds. Some nickel mayco-precipitate with the iron resulting in losses of the pay metal and inaddition the neutralizing agent represents an additional operating costof the process. A further advantage of the use of higher temperature isan improvement in the solid/liquid separation properties of the finalleach slurry, with higher settling rates being achieved with a higherleaching temperature.

In some instances, it may be advantageous to employ both an atmosphericpressure leaching stage and a subsequent moderate pressure leachingstage, in series. Sulfur dioxide, or another reductant, for the leachingof oxidized manganese minerals and contained cobalt and nickel values ismost conveniently added during the atmospheric leaching stage. Asubsequent pressure leaching stage can still be used to achieve theadvantages noted in the preceding paragraph.

Nickel extractions of at least 80% and as high as 90-95%, withcorresponding iron extractions as low as 5-10%, are obtained with thecurrent invention. In addition, the leach residue typically contains nomore than about 1-2% of sulfur, indicating that the iron compounds inthe residue are primarily not of the jarosite type.

This leaching process of the present invention, in the case wheresulfuric acid is used as the lixiviant, differs from the prior art, forexample U.S. Pat. No. 3,244,513, in that the latter employs a roastingstep after acid blending in order to convert any iron sulfate formedduring acid blending to iron oxide. Sulfur trioxide (SO₃) gas isreleased during this conversion and this reacts with non-ferrous metaloxides, e.g. NiO and MgO, to form non-ferrous sulfate salts. Thus, thenon-ferrous metals are preferentially sulfated in the process of theprior art.

However, is has surprisingly now been found that the roasting step,which is likely to be capital cost intensive and require substantial,expensive fuel energy for achieving the desired roasting temperatures,e.g. 500-700° C., is not necessary to achieve selective sulfation of thenon-ferrous metals, thus resulting in a pregnant leach solution afterwater leaching of the ore/acid blend, which contains most of the solublenon-ferrous metals, particularly nickel and cobalt, and relativelylittle of the iron. All that is required is to provide sufficient timeand temperature during water leaching to effect the desired reactions.Elimination of the roasting step thus represents a major advantage overthe prior art.

After water leaching, the leach slurry is subjected to solid/liquidseparation by filtration or thickening to produce a pregnant leachsolution containing most of the nickel and cobalt contained in the oreand a solid residue containing most of the iron in the ore.Advantageously, thickening is carried out in a series of thickeners withcounter-current flow of a wash water stream and the leach slurry inorder to wash most of the entrained metal values out of the leachresidue, a method called counter-current decantation (CCD). The metalvalues report preponderantly to the thickener overflow of the firstthickener, which is the pregnant leach solution.

The pregnant leach solution proceeds to nickel and cobalt recovery bymethods known to those skilled in the art, such as solvent extraction,ion exchange, sulfide precipitation using sulfiding agents, e.g.hydrogen sulfide, or hydroxide precipitation, using for example magnesiaas the precipitating agent.

The nickel and cobalt can also be recovered from the leach slurrywithout prior solid/liquid separation, using the resin-in-pulp process.In this process, an ion exchange resin which extracts nickel andpossibly cobalt is added directly to the leach slurry. After theextraction is complete, the resin is separated from the nickel-depletedleach slurry by screening. After washing the resin to remove solids, thenickel can be eluted from the resin with a fresh acid solution.

Prior to or during nickel and cobalt recovery by one of theaforementioned methods, the leach solution may be neutralized with abase, such as calcium carbonate, magnesium oxide, sodium carbonate orthe like, to neutralize free acidity remaining from the leach processand precipitate small amounts of ferric iron, aluminum, and chromium,while minimizing co-precipitation of nickel and cobalt. This process maybe carried out in a single or multiple steps separated by intermediatesolid/liquid separations.

In one embodiment of the invention, the first stage of neutralizationmay be carried out prior to separating the leach residue from the leachsolution. The combined leach and neutralization residue may then beseparated from the partially neutralized leach solution by filtration orthickening, as described above. A second stage of neutralization maythen still be desirable, depending on the method selected for nickel andcobalt recovery from the pregnant leach solution. After this secondstage of neutralization, the resulting neutralization residue may beseparated from the neutralized leach solution by filtration orthickening. This second-stage neutralization residue is ideally returnedto the first stage neutralization to re-dissolve any co-precipitatednickel and cobalt.

It has surprisingly been found that it is not necessary to separate thelimonite from the saprolite in the laterite ore in order to treat themseparately and distinctly in the leach process. Further, the process isnot limited to any particular saprolite/limonite ratio, as long as atleast a certain minimum saprolite/limonite ratio is met. The quantity ofacid added to the ore may be adjusted for a range of saprolite/limoniteratios based on the non-ferrous metals content of the saprolite/limoniteore mixture.

It has surprisingly also been found that it is not necessary to add anyiron precipitating agents to the leach slurry, as in U.S. Pat. No.3,793,432. Addition of iron precipitating agents such as sodium,potassium and ammonium ions is a disadvantage in that they promote theformation of iron-bearing jarosite compounds, which arethermodynamically unstable, slowly decomposing over time and releasingsulfuric acid, which can cause the re-dissolution of metal impuritiespresent in the leach residue and leading potentially to contamination ofthe environment. In addition, extra sulfuric acid is required duringleaching to satisfy the requirement for sulfate ions in the formation ofjarosite.

In the process of the present invention, where sulfuric acid, forexample, is added to the ore, nickel, cobalt and iron, as well as othernon-ferrous metals in the ore, are unselectively sulfated, that is,converted to sulfate salts, during the acid/ore blending step. Thesulfates dissolve easily in water during the subsequent leach step.However, nickel and cobalt extraction are incomplete initially becausethe quantity of acid added initially is not sufficient to convert all ofthe metals, including iron, to the sulfate salts. Thus, considerableiron goes into solution along with nickel, cobalt, magnesium, aluminum,chromium, copper and zinc. However, by continuing the leach processafter the initial sulfate salt dissolution phase, the iron content ofthe solution decreases and further nickel and cobalt are extracted fromthe solid phase. The exact iron leaching and re-precipitation reactionsare unknown, but the sulfur content of the final residue is low, usuallyless than 2%, indicating that jarosite does not form to a significantextent.

In one preferred embodiment of the invention, an iron-bearing “seed”material is added to the leach slurry at the start of the water leach,in order to accelerate the precipitation of dissolved ferric iron andthe extraction of remaining nickel and cobalt from the solid phases. Thesurfaces of the seed particles provide low-activation energy sites forhydrolysis and precipitation of iron, for example as ferric hydroxide,goethite, or hematite. This seed material is ideally a portion of thefinal leach residue itself, which contains precipitated iron compounds.One method of carrying out the process with seed recycle is shown inFIG. 2.

In a further embodiment of the invention, the limonite portion of theore is first blended with the concentrated acid, the amount of acidbeing calculated in the same manner as discussed previously, followingwhich the saprolite portion of the ore is added to the limonite/acidmixture. The saprolite ore may be added to the limonite/acid mixtureeither as crushed saprolite or as ground saprolite. In the former casethe ore/acid mixture is ground prior to water leaching. The grinding maybe carried out without the addition of water. Advantageously, water isadded to the ore/acid mixture and wet grinding is employed prior toatmospheric water leaching. In the latter case, the saprolite ore can beground in the dry condition and added to the acid/limonite mixture alongwith water, or ground wet and added to the limonite/acid mixture as aslurry or filtercake. The resulting blended ore and acid is then waterleached as described previously. This method provides the maximumopportunity to sulfate the goethite component of the laterite ore. Aflowsheet for carrying out this embodiment of the invention is shown inFIG. 3.

The following examples illustrate the method of the present invention.The ore used in these examples came from a Central American lateritedeposit. The limonite and saprolite fractions of the ore used inExamples 1-3 had the compositions given in Table 1. The saprolite orewas crushed to approximately −6.4 mm before use in the tests. TABLE 1 %Ni % Fe % Mg % Moisture Limonite Ore 1.41 47.7 0.67 34.5 Saprolite Ore3.17 8.7 17.8 21.3

EXAMPLE 1

Approximately 1 kg of saprolite ore was wet ground in a ball mill toapprox. 100% passing 100 mesh. The ground slurry was filtered to produceground saprolite at 41% moisture content. 425.2 g of the moist groundsaprolite was blended with 381.7 g of moist limonite ore to produce 500g of leach feed material with a limonite/saprolite ratio of 1:1 on a drybasis.

The ore mixture was placed in a 4.5 liter narrow-necked glass bottle.The bottle was rolled at an angle inclined slightly to the horizontal atapproximately 47 to 48 rpm on a bottle rolling device. 312.5 g of 96%sulfuric acid was added to the ore mixture in the bottle over a periodof about 30 minutes. The ore and acid were blended for approximately 30min after all of the acid had been added. At the completion of blendingthe ore and acid had formed a semi-fluid mass and the temperature hadrisen to between about 70 and 100° C.

The bottle was removed from the rolling device and the blend of acid andore was allowed to cure at ambient temperature for approx. 72 hours. 622mL of water were then added to the cured mass and the mixture wasstirred until a uniform leach slurry was formed. The leach slurry wastransferred to a 2-liter cylindrical reaction kettle equipped with amechanical stirrer, 4 plastic baffles, and a tight-fitting lid equippedwith a water condenser open to the atmosphere. The reaction kettle washeated by an external, electrical heating mantle.

The leach slurry was heated to and held at a temperature ofapproximately 96-99° C. for a period of 48 hours while stirringvigorously. After 5 hours of leaching, 129 g of finely ground, technicalgrade hematite was added to the reactor to act as a “seed” for theprecipitation of iron. Samples of the leach liquor were takenperiodically for chemical analysis. At the end of the 48-hour leachingperiod, the entire slurry was filtered. The filtercake was repulpedtwice in fresh water to wash out the entrained metal values. The cakewas then dried and weighed. The dry solids, filtrate and the combinedwash water were assayed separately.

The results of this test are given in Tables 2 and 3 below. TABLE 2 Time(h) [Ni] g/L [Fe] g/L [Mg] g/L [H₂SO₄] g/L 1 8.8 70 27 <0.5 2 8.0 49 27<0.5 5 9.5 24 35 <0.5 24 8.7 14 35 8 32 8.3 12 32 9 48 8.8 10 28 16

TABLE 3 Ni (% or g/L) Fe (% or g/L) Mg (% or g/L) S (%) Final Solution9.3 12.8 39 — Final Residue 0.34 42.9 0.52 0.98 Calculated 85.2 11.794.9 — Extractions**Ni, Fe and Mg extractions based on final residue and solution weights,volumes and assays.

The results show that a high extraction of nickel was achieved whileextracting very little of the iron present in the ore. In addition, thesulfur content of the residue was quite low and the Fe/S ratio wasapproximately 44:1. Less than 5% of the sulfate contained in thesulfuric acid deported to the residue. If all of the iron in the ore hadformed jarosite, the Fe/S ratio in the final residue would have beenapproximately 4.3. In addition, X-Ray diffraction analysis of theresidue indicated the presence of only hematite, goethite and a spinelphase originally present in the ore.

The kinetic data indicate that slightly more than 24-32 hours ofleaching was required to approach the terminal iron concentration in thesolution. The residual iron could be removed from solution byneutralization with a base, for example, limestone, without siqnificantlosses of nickel to the solid phase.

EXAMPLE 2

A sample of 381.7 g of limonite ore was blended with 317.7 g of crushedsaprolite ore and 95 g of water in a 2-liter flat-bottomed glass beaker.The water was added to simulate the expected moisture content ofrun-of-mine ore, recognizing that the samples had dried somewhatcompared to their in-situ condition. 312.5 g of 96% sulfuric acid wasadded to the beaker over approximately 30 minutes. The acid was blendedwith the ore using a stirrer rotating at about 60 rpm. The acid additionwas sufficient to give an acid to ore ratio of about 600 kg H₂SO₄ pertonne of ore (dry basis).

The ore and acid formed a semi-fluid mass which was poured into ashallow pan for curing at ambient temperature for approximately 72hours. Some of the mixture was not recovered during this operation. Itwas estimated that about 13% of the mixture was not recovered, based onweight recovery.

After this period of curing, the acid/ore mixture, which had hardenedconsiderably, was broken into pieces and transferred to a small grindingmill. 300 g of water were added to the mill and the mixture was groundwith stone media for approximately 1 hour to reduce the maximum particlesize to about 100 mesh. Additional water was added to the mill to washthe slurry to a 2-litre leach reactor. 1,858 g of leach slurry wasprepared in this way. The reactor was heated to a temperature of 95 to105° C., while stirring, and leaching was allowed to continue for 44hours. After 5 hours of leaching, 129 g of finely ground, technicalgrade hematite were added to the reactor to act as a “seed” for theprecipitation of iron. Samples of the leach liquor were takenperiodically for chemical analysis. At the end of the 48-hour leachingperiod, the entire slurry was filtered. The filtercake was repulpedtwice in fresh water to wash out the entrained metal values. The cakewas then dried and weighed. The dry solids, filtrate and the combinedfresh water were assayed separately.

The results of this test are given in Tables 4 and 5 below. TABLE 4 Time(h) [Ni] g/L [Fe] g/L [H₂SO₄] g/L 1 4.6 25.6 <0.5 2 5.36 29.7 <0.5 56.15 15.5 5 24 4.73 4.74 12 44 5.49 7.55 16

TABLE 5 Ni (% or g/L) Fe (% or g/L) Mg (% or g/L) S (%) Final Solution5.51 8.7 22.0 — Final Residue 0.41 42.8 0.47 0.83 Calculated 81.1 11.394.7 — Extractions**Ni, Fe and Mg extractions based on final residue and solution weights,volumes and assays.

This test simulated one of the preferred embodiments of the currentinvention, in that saprolite ore was not ground before blending withlimonite and adding acid. Instead, the acid/ore mixture was ground aftercuring.

The results of this test were similar to those of Example 1, althoughnickel extraction was slightly lower.

EXAMPLE 3

This test was carried out in a fashion similar to that of Example 2,with the following exceptions. The amounts of limonite ore, crushedsaprolite ore, water, acid, and hematite seed used during the acidblending and subsequent leaching process were 336.9 g, 280.3 g, 84 g,275.8 g, and 113 g, respectively. The proportions of ore, water and acidwere the same as in Example 2.

After blending the ore with acid, the mixture was transferred to ashallow pan and allowed to cure for only one hour before transfer to agrinding mill for wet grinding and subsequent water leaching. After onehour of curing the mixture was still fluid and could be poured into themill. 1,776 g of leach slurry were prepared and leached exactly as inExample 2. Recovery of the acid/ore mixture was approximately 96%.

The results of this test are given in Tables 6 and 7 below. TABLE 6 Time(h) [Ni] g/L [Fe] g/L [H₂SO₄] g/L 1 5.53 22.7 <0.5 2 6.37 24.2 <0.5 57.59 16.7 4 24 7.4 9.14 13 44 7.12 10.8 16

TABLE 7 Ni (% or g/L) Fe (% or g/L) Mg (% or g/L) S (%) Final Solution6.97 10.2 34.0 — Final Residue 0.37 43.0 0.45 0.81 Calculated 84.2 10.895.2 — Extractions**Ni, Fe and Mg extractions based on final residue and solution weights,volumes and assays.

The results of this test, which are similar to the previous examples,demonstrate that it is not necessary to cure the ore/acid mixture for along period of time prior to leaching.

EXAMPLE 4

The ore used in this test had the compositions given in Table 8. TABLE 8% Ni % Co % Fe % Mg % Si % Moisture Limonite 1.31 0.2 47.2 0.63 2.6741.9 Ore Saprolite 3.13 0.01 6.0 20.0 18.8 14.4 Ore

409.6 g (wet basis) of the limonite ore was placed in a small porcelainball mill charged with 0.6-1.9 cm porcelain balls. Prior to adding theore, the mill and grinding media were preheated to approx. 100° C. Thiswas done to simulate the thermal conditions that would be present if theprocess were carried out on a continuous basis. The mill was fitted witha plastic lid with a hole in its centre through which sulfuric acidcould be added. The mill was rotated on the same rollers used inExample 1. 338.5 g of 96% sulfuric acid were added to the mill over aperiod of approximately 1.5 minutes and the limonite ore and acid wereallowed to react for a period of 15 minutes. The temperature inside themill was monitored with a hand-held laser thermometer during thereaction period. The temperature measured varied from 97 to 121° C.,with the highest temperature recorded about 1 minute after the acidaddition was completed. At the completion of the reaction period, themill was removed from the rollers, the lid was removed and 306.1 g (wetbasis) of crushed saprolite ore was added to the limonite/acid mixturealong with 510 g of a 30 g/L Mg (as MgSO₄) aqueous solution. MgSO₄solution was used to simulate the effect of using the barren solutionremaining after purification and nickel/cobalt recovery as make-up waterin the process. A tight-fitting porcelain lid was used to close the milland the total ore/acid mixture was ground for approximately 60 minutes,until most of the solid particles present were less than 100 mesh insize.

Based on the weights and moisture contents of the ore samples, theoverall saprolite/limonite ratio was 1.1 on a dry basis and the overallacid/ore ratio was 0.65 (based on dry ore and 100% H₂SO₄).

The contents of the mill were then discharged into a reaction kettleidentical to that used in Example 1. A coarse screen was used to capturethe grinding media and 764 g of the 30 g/L Mg solution was used to rinsethe ball mill and the grinding media. The rinse solution was added tothe reaction kettle.

The reaction kettle was heated to 95-100° C. while stirring. Leachingwas carried out for 24 hours. During the first 5 hours of leaching,sulfur dioxide gas was bubbled into the leach slurry to control theoxidation reduction potential of the slurry at between 540 and 600 mV(versus a saturated Ag/AgCl reference electrode). Samples of slurry weretaken throughout the leaching period, filtered, the solids washed, andthe solution and solids assayed.

At the completion of the leaching period, 20% limestone slurry was addedto the leach slurry over approximately 2 hours to neutralize the leachslurry to a pH of 3.0 at 95° C. The neutralized leach slurry wassampled, then the entire slurry was filtered.

The results of the leaching period are given in Table 9. TABLE 9 Time(h) [Ni] g/L [Co] g/L [Fe] g/L [H₂SO₄] g/L 1 4.92 0.229 60.8 <0.5 2 5.330.248 50.0 <0.5 4 6.22 0.290 33.9 <0.5 5 6.52 0.303 26.5 <0.5 24 7.360.34 14.4 5.0

The solids at the completion of the leaching period assayed 0.21% Ni,0.006% Co, 31.3% Fe, 0.65% Mg, 14.9% Si and 1.39% S. The Ni and Coextractions, calculated using a “silicon tie” (assuming zero siliconleaching, calculating leach residue weight using silicon assays of oreand residue, and using ore and residue weights and assays to calculateextractions), were approximately 93.1% and 95.5%, respectively. The ironextraction was roughly 9%.

After neutralization to pH 3, the solution assayed 6.23 g/L Ni, 0.29 g/LCo, and 3.24 g/L Fe. The solids assayed 0.22% Ni, 0.009% Co, 26.5% Feand 11.7% Si. Approximately 74% of the iron in the leach solution wasprecipitated during neutralization; thus net iron extraction afterneutralization was only about 2%. Approximately 2% of the nickel and 4%of the cobalt were co-precipitated with iron during neutralization.

This example illustrates several key features of one of the preferredembodiments of the invention: high nickel extraction can be achieved byfirst reacting sulfuric acid directly with the limonite ore and thenadding the saprolite ore; the addition of a reductant to the leachslurry, in this case sulfur dioxide gas, enhances cobalt extraction;there is no need to cure the acid/ore reaction mixture to achieve highextractions; the reaction between sulfuric acid and ore is extremelyrapid and extremely exothermic, such that very high temperatures can beachieved during the acid/ore reaction, thereby enhancing the reactionkinetics and permitting the equipment required to carry out thesulfation step to be very compact in size; and partial neutralization ofthe leach slurry following leaching can be used to eliminate most of thesmall proportion of iron which leaches with minimal nickel and cobaltlosses.

It will of course be appreciated by those skilled in the art that manyvariations of the process would be possible within the broad scope ofthe present invention. Those skilled in the art will appreciate that theinvention upon which the description is based may be utilized in otherembodiments that carry out the purposes and fulfill the objects of thepresent invention.

The process of the current invention is widely applicable to nickellaterite ore bodies containing both limonite and saprolite, whichincludes the majority of such ore bodies. The saprolite/limonite ratioin the process may vary widely as long as a certain minimumsaprolite/limonite ratio is used in the process. When using sulfuricacid, for example, the minimum ratio is determined approximately bycalculating the amount of non-ferrous components of the saprolite orewhich can be sulfated by the sulfuric acid that would be generated byhyudrolyzing and precipitating the “sulfatable” iron content of thelimonite and saprolite ore as an oxide or hydroxide of ferric iron. Theratio may be any value equal to or greater than this minimum, providingthat the acid addition is calculated as described previously, that is,sufficient acid to react with the sulfatable, non-ferrous components ofthe linionite and saprolite ore, plus a slight excess to account for asmall amount of soluble iron and to drive the reactions to completion.

The above disclosure is intended to be illustrative while the scope ofthe invention is defined by the following claims.

1. A process for leaching laterite ores containing limonite andsaprolite, comprising the following steps: (a) mixing the limonite andsaprolite with sufficient concentrated mineral acid to form salts withsoluble non-ferrous components of the ore; (b) water leaching themixture of step (a) at a temperature and for a time sufficient to causehydrolysis of dissolved ferric iron whereby an iron containingprecipitate is formed, while substantially dissolving at least one ofeither nickel or cobalt into the leach solution; and (c) recovering atleast one of either nickel or cobalt compounds from the leach solution.2. A process as recited in claim 1, in which step (a) is carried out byfirst mixing the acid with the limonite, and subsequently adding thesaprolite.
 3. A process as recited in claims 1 or 2, in which at least aportion of the ore is crushed to promote intimate mixing of the acidwith the ore.
 4. A process as recited in claim 2, in which the saproliteis crushed separately and then added to the mixture of limonite andacid.
 5. A process as recited in claim 1, in which combined limonite andsaprolite is crushed and ground prior to mixing with the acid.
 6. Aprocess as recited in claims 1, 2, 4 or 5, in which the water leachingof step (b) is carried out in an autoclave at a temperature up toapproximately 150° C.
 7. A process as recited in claim 6, in which thewater leaching of step (b) is carried out at a temperature range ofapproximately 95-105° C.
 8. A process as recited in claims 1, 2, 4 or 5,in which the water leaching of step (b) is carried out at a temperaturerange of approximately 95-105° C. and at a pressure range ofapproximately 15-70 psia.
 9. A process as recited in claims 4 or 5, inwhich the water leaching of step (b) is conducted in two stages, thefirst stage being carried out at atmospheric pressure and at atemperature up to the boiling point of the leach solution, and thesecond stage being carried out in an autoclave at a temperature up toapproximately 150° C.
 10. A process as recited in claims 4 or 5, inwhich the acid is selected from the group of sulfuric, hydrochloric, andnitric acid, or mixtures thereof.
 11. A process as recited in claim 8,in which the acid is sulfuric acid and the salts formed in step (a)include sulfates.
 12. A process as recited in claims 1, 2, 4 or 5, inwhich the recovering of at least one of either nickel or cobaltcompounds from the leach solution in step (c) includes adding an ionexchange resin to the leach solution without prior solid/liquidseparation.
 13. A process as recited in claims 1, 2, 4 or 5, in whichthe leach solution is first separated from the precipitate prior to therecovering of at least one of either nickel or cobalt compounds from theleach solution in step (c).
 14. A process as recited in claims 1, 2, 4or 5, in which a reductant is added during the water leaching of step(b) to enhance the dissolution of cobalt from the ore.
 15. A process asrecited in claim 14, in which the reductant is selected from the groupof sulfur dioxide, hydrogen sulfide, soluble bisulfite and sulfitecompounds, or soluble ferrous iron compounds.
 16. A process as recitedin claim 1 or 2, further comprising the step of curing the mixture ofstep (a) prior to the water leaching of step (b).